Progress in lead anode mud treatment technology

The progress of lead anode mud treatment technology is mainly towards the development of wet process. Its focus is to reduce the environmental pollution caused by arsenic and lead, improve the direct yield of gold and silver , and directly obtain finished products without electrolysis, further shortening the production cycle.

A three ferric chloride leaching process

The characteristics of the ferric chloride leaching process proposed by Shenyang Smelter are: after the lead anode mud is leached with ferric chloride, copper , bismuth , antimony , etc., ammonia leaching and silver extraction, leaching slag smelting and electrolysis, the process flow is shown in Figure 1. . The lead and bismuth mud components used in the test are shown in Table 1.

Figure 1 Process of leaching lead anode mud from ferric chloride

Table 1 Lead anode mud composition (%)

(1) Treatment of leaching and leachate

The results showed that:

1. As the ratio of iron to iron decreases, the amount of iron trichloride increases, and the leaching rates of As, Sb, Bi, Cu and other metals increase, that is, Pb fluctuates around 50%. When the ratio of feed to iron is 1:0.72~0.76 (0.74 is equivalent to 140g/Lfe ³⁺ );

2. Acidity (excluding the acidity of FeCl ³ solution) is in the range of 0.4-0.6mol/L, and As, Sb, Bi, Cu, etc. have higher leaching rate, but when the acid is low, the filtration speed is slower, to 0.5. Mol/L is suitable;

3, the solid-liquid ratio is 1:5 ~ 7, generally used 1:4.5 ~ 5.5, stirring for 2h, the leaching rate of As, Sb, Bi, Cu are relatively high;

4. The leaching rate of each metal increases with the increase of temperature, and increases greatly between 50-55 ° C and 60-65 ° C. The change of heating temperature is not obvious. Considering the energy consumption and equipment corrosion, the general choice is 60. ~65 ° C. Under the selected conditions, the leaching results are shown in Table 2.

Table 2 FeCl3 leaching solution, slag composition

Hydrolysis: The purpose of hydrolysis is to separate ruthenium from other metals. According to the theory, the hydrolysis of zinc can be carried out under the condition of high acidity, as long as the concentration of chloride ions in the hydrolyzate is reduced to a certain limit, even higher acidity. The hydrolysis of hydrazine is also thorough, while other metals such as copper and strontium are not hydrolyzed. The silver in the solution is also precipitated together with the cerium and is enriched in SbOCl in the cerium oxide. Therefore, the leachate is diluted with water, and SbCl _{3 is} hydrolyzed, and the reaction is:

SbCl ₃ +H ₂ O→SbOCl↓+2HCl

Silver precipitated with AgCl. The precipitation rate of bismuth and silver is more than 99%. Cu and Bi remain in the solution. The effect of water dilution ratio on the precipitation rate of bismuth and silver is shown in Table 3. The table shows 6 parts. According to the analysis of the hydrolyzate, é“‹ coprecipitated about 5%.

Table 3 Effect of water dilution ratio on precipitation of Cu and Bi

Neutralization: After hydrolysis and sedimentation, the pH is about 0.5, and it is neutralized with sodium carbonate or calcium carbonate to a pH of 2.0 to 2.5, and the ruthenium can be completely precipitated and recovered. A small amount of silver remaining in the hydrolysis also precipitated together. Copper is still left in the solution. If it is ensured that there is not too much Fe ^{3+ in the} leaching solution, the quality of the precipitate is very high. At this time, the copper and copper have been separated. The effect of neutralization pH on cerium precipitation is shown in Table 4.

Table 4 and the effect of pH on precipitation enthalpy

The neutralized stock solution contained Sb 45.5 g/L, Bi 29.2 g/L, Cu 8.47 g/L, Pb 5.35 g/L, and was neutralized with 20% Na ₂ CO ₃ after hydrolysis.

After neutralization and sinking, the solution contains Cu2.3g/L, which can be treated by sodium sulfide precipitation or iron filings- limestone neutralization. When using Na ₂ S, the temperature was 30 ° C, stirring for 1 h, the amount of Na ₂ S was 120% of the amount of copper, and the liquid composition after precipitation was: Pb 0.0013 g / L, Cu < 0.001 g / L, Sb 0.016 g / L,

Bi0.0019g / L, basically meet the waste discharge standards. The latter replaces copper with iron filings to recover copper into sponge copper. After replacement, the liquid composition is:

Pb0.0013g/L, Cu0.001g/L, Sb0.22g/L, Bi0.23g/L, As0.006g/L, neutralized with limestone to pH=8～9, waste liquid component: Pb0.001g /L, Cu <0.001g / L, Sb0.003g / L, Bi 0.001g / L, As trace, to achieve direct discharge standards. Accordingly, it is recommended to use the iron filings-limestone method to obtain pure sponge copper, and the cost is less.

(2) Recovery of silver

The yang grade mud is leached by FeCl3. More than 95% of the silver and all the gold are enriched in the leaching slag of ferric chloride. The slag contains more than 50% of silver and can be treated by mature smelting electrolysis. Such as soda, carbon powder (about 3%) smelting, crude silver direct yield 95% ~ 97%. The silver powder obtained by silver electrolysis is ingot by the ingot, and the gold enters the anode mud, and after the nitric acid is boiled to silver, the finished gold is treated by electrolytic refining or chemical treatment. For example, 783.3g is charged, 778g is silver, 96.79% is silver, and the crude silver yield is 96.14%. The crude silver gold was 0.081%, the gold was 778×0.08%=0.63 g, the raw material input gold was 0.66 g, and the direct yield of gold was 95.4%. It can also be treated by wet method, that is, with ammonia solution, liquid-solid ratio of 5:1, temperature of 50-70 ° C, and the leaching solution is reduced by hydration with ammonia, and the recovery rate of silver is >99%. The ammonia leaching residue is reduced and smelted into crude silver electrolysis, and gold is recovered from the anode slime.

Second, the full wet process - HCl + NaCl leaching

The ruthenium and osmium in the lead anode slime were separated by leaching with HCl+NaCl and pre-separated and recovered, and then the gold and platinum , palladium , sulfite reduction gold and iron powder were used to replace the platinum and palladium concentrate in the sulfuric acid medium; The gold slag is extracted by ammonia leaching, and the hydrated hydrazine is reduced. The process flow is shown in Figure 2.

Figure 2 Lead anode mud wet processing process

The anode mud composition for the test is: Au 0.4% to 0.9%, Ag 8% to 12, Sb 40% to 45%, Pb 10% to 15%, Cu 4% to 8%, As 0.87%, Fe 0.62%, Zn0. 03%, Sn 0.001%.

When the liquid-solid ratio is 6,70-80 ° C, hydrochloric acid 1.5 mol / L, [Cl -] = 5 mol / L, stirring leaching for 3 h, the leaching rate is Sb99%, Pb29% ~ 53%, Bi98%, Cu90%, As90%.

The conditions of chlorination are: liquid-solid ratio 6, H ₂ SO ₄ 100g / L, NaCl 80g / L, 80 ~ 90 ° C, the amount of NaClO ₃ is 3.5% ~ 5% of the weight of the anode mud, chlorination 2h, gold leaching rate >99.5%, Na ₂ SO ₃ reduced the grade of 95% ~ 98% gold powder, gold yield > 98%.

Ammonia leaching silver: liquid-solid ratio 5-8, 1:1 ammonia water, 30 ° C, stirring leaching for 2 h, silver leaching rate of 99.5%, hydrazine hydrate reduced silver powder, silver direct yield of 97%.

The process can also comprehensively recover other valuable metals such as lead, antimony, bismuth and copper, and the direct yields are: 84%, 70%, 85%, 92%.

Third, chlorine salt leaching - silicon fluoride de-lead - melt casting alloy - gold and silver electrolysis process

The scientific research institute of Shuikoushan Mining Bureau passed the experiment on the composition of lead anode mud, which is shown in Table 5, with low gold content and relatively high content of bismuth, antimony, copper and lead. The experimental results suggest that the chloride salt leaching-conversion-silicofluoric acid delead-de-lead slag directly melt-casts gold-silver alloy-gold-silver electrolysis, which can effectively recover precious metal gold and silver and comprehensively recover valuable metals such as copper, lead, antimony and bismuth. The purpose is to realize the smokeless gold and silver process.

Table 5 Chemical composition of lead anode mud (%)

(1) Chloride salt leaching

Liquid-solid ratio = 4 ~ 5: 1, [HCl] network = 2.5mol / L, [Cl ^- ] = 5mol / L (the amount of salt added is 30% ~ 40% of the anode mud); metal leaching rate Cu, Sb, Bi ≥ 98%, Ag 1%, and Au does not substantially leaching. The composition of the leachate is: Au<0.001g/L, Ag0.15g/l, Cu9.56g/L,

Sb32.47g/L, Bi24.92g/L. The leaching slag composition was: Au841g/t, Ag31.52%, Cu0.46%, Sb0.041%, Bi0.013%.

(2) Conversion and restoration

After the lead anode mud is leached by the chloride salt, Ag and Pb are substantially insoluble in the general medium by AgCl, PbCl2 and PbSO4, so PbCl _{2 is} converted into PbCO ₃ . (ie, PhCl ₂ + Na ₂ CO ₃ → 2NaCl + PbCO ₃ ), the conversion conditions are: alkali-to-lead ratio = 1.2:1, liquid-solid ratio = 7:1, temperature 80 ° C, time 3 h.

(3) Silicic acid leaching

Silicic acid (nitrogen and acetic acid can react with PbCO3) leaching the conversion slag, the purpose is to separate the PbCO3 form into Au and Ag.

The use of silicic acid is considered to contain a large amount of AgCl in the conversion slag. In order to avoid the loss of silver and the regeneration of silicic acid, the leaching of lead acid by fluoric acid is selected. The reaction is:

PbCO ₃ +H ₂ SiF ₆ =PbSiF ₆ +H ₂ O+CO ₂ ↑

PbSiF ₆ +H ₂ SO ₄ =H ₂ SIF ₆ +PbSO ₄ ↓

The chlorofluoric acid leaching conditions are: liquid-solid ratio = 4 ~ 5: 1, temperature 30 ° C, silicic acid concentration > 12%, silicic acid leaching lead removal rate of 98% or more. The slag contains less than 1% lead.

The lead-free slag can directly melt the alloy and the recovery rate of gold and silver.

Fourth, wet method comprehensive recovery of lead anode mud

The composition of lead anode mud of a factory in Hunan is: gold 65.6g/t, silver 21.0%, lead 29.67%, bismuth 12.35%, strontium 11.71%, arsenic 2.57%, copper 1.01%. In the HCl+NaCl gold wet process, under the condition of 70-80 °C, using NaClO ₃ as oxidant, the leaching potential is controlled at 400-420 mV, and Sb, Bi, Cu and As are leached to ensure the residual acid of the leaching solution is higher than 1.5 mol. /L, hydrolyzed and recovered from the leachate, replaced with iron scraps, neutralized with waste liquid, and treated with arsenic removal. The HCl+NaCl process is simplified. Directly separate the base metal.

Chlorinated gold: bismuth metal leaching residue, in the H ₂ SO ₄ + NaCl system, at 80 ~ 90 ° C, with NaClO ₃ as an oxidant, to ensure a leaching potential of 1200mV. Fully chlorinated material (3h), the leaching residue contains gold ≤15g/t, the leaching solution adsorbs gold with activated carbon, and the absorption rate is 99.29%.

Ammonia leaching silver: chlorinated gold slag with 1:1 ammonia water, liquid-solid ratio = 10:1, leaching for 2h, silver output rate is greater than 96%, the precious liquid is reduced by hydrazine hydrate at 50-60 ° C to obtain a purity of 99.7 % sponge silver, the reduction rate is 99.9%, and the total recovery rate of gold and silver in the whole process is ≥99%.

After the ammonia is immersed in silver, the slag is melted into a lead ingot reverse electrolysis system.

5. Liquid chlorine leaching lead anode mud

The lead anode mud of a certain plant is: Pb 10% to 15%, Sn 15% to 20%, Bi 3% to 5%, As 15% to 20, Sb 15% to 25%, Ag1% to 1.5%, and Au 5 to 30 g/t. Through the test, a liquid chlorine leaching-extraction treatment process was developed, as shown in FIG. The leaching operation was carried out in a 2 m ³ enamel reaction tank. Add 5.5mol/L hydrochloric acid solution, steam indirectly to 40~50°C, gradually add lead anode mud (liquid-solid ratio 4:1) under mechanical stirring, and start chlorine gas leaching after 0.5h addition. As the leaching releases heat, the mine will raise the temperature to 80-90 °C. Therefore, the starting liquid temperature must be strictly controlled to avoid causing the slurry to boil. During the leaching process, As, Sb, Sn, Bi, Pb, Ag, etc. are all converted into corresponding arsenic acid and chloride. Most of the lead chloride produced remains in the slag and a small portion enters the solution. When cooling, most of the lead chloride crystals precipitated, and the lead content of the solution after cooling and crystallization can be reduced to 1-2 g/L.

Figure 3 Chloride leaching of lead anode mud liquid - extraction process

The chlorine gas is leached for 1 to 2 hours until the slag is grayish white, and the sedimentation is fast. When the supernatant liquid is dark brown and is not turbid, the chlorine is stopped. Then 2.5% of the raw anode mud was added and stirred for 1 h to remove excess free chlorine and to reduce precipitation of the precious metal entering the solution. At this time, the pentavalent anthracene is also reduced to trivalent to prevent the high-priced crucible from destroying the organic phase in the next extraction. The leaching sulphur slurry is clarified, and the supernatant liquid is sent to a storage tank for extraction and separation. The tin is first extracted with P350, and then extracted with N235. The raffinate is evaporated and then caustic soda is neutralized to recover arsenic and antimony. The intermediate products produced are sent separately for purification.

The liquid chlorination method mainly contains lead chloride and silver chloride, and the silver and lead are separated by silver chloride immersion:

AgCl+2NH ₄ Cl→Ag(NH ₃ ) ₂ Cl+2HCl

When leaching, ammonia gas is introduced into the hydrochloric acid solution to maintain the solution pH=9 or control free ammonia about 35g/L. After the solid-liquid separation, the silver in the ammonia immersion liquid is reduced by hydrazine hydrate. Silver mud washing, drying and ingot casting. After the reduction of the precipitated silver, the liquid is returned to the secondary leaching of the leaching slag, and the secondary leaching solution is used for the primary leaching of the liquid chlorine leaching slag.

After the liquid chlorine leaching residue is immersed in silver by the second ammonia, the lead is immersed in salt. The salt immersion solution is diluted with water and neutralized with sodium carbonate to recover lead carbonate. The precipitate was dissolved in acetic acid to prepare crystal lead acetate. The salt leaching lead slag still contains a small amount of gold, silver and non-ferrous metals, and is recovered by other methods.

6. Caustic soda leaching

In the lead anode mud, As, Sn, Pb Sb, and Te are all present as oxides, and when the caustic soda is leached, the corresponding sodium salt can be transferred into the solution. Gold, silver, copper, bismuth, etc. in the lead anode mud remain in the leaching residue. Slag smelting can reduce pollution and basically eliminate lead. Practice has shown that when the lead anode mud is oxidized to a grayish white color after long-term storage, the texture is loose and the leaching separation effect is better.

A plant has conducted a caustic soda leaching test on four lead anode muds of different components after long-term storage of oxidation. The lead anode mud is ground and mixed in a ball mill at a normal temperature and liquid-solid ratio of 3:1, and ground to 60 mesh. The liquid-solid ratio is 10:1 in the iron drum stirred tank, and the initial concentration of caustic soda is 180-200 g/ L, stirring at a temperature of 95 to 100 ° C for 2 h. The leaching ratio of each component was As97%, Sn94%, Pb90% Sb 70% to 98%, and Te0 to 40%. The yield of leaching residue is 8% to 40%, and gold, silver, copper and strontium are all retained in the slag, and the enrichment ratio is 2.5 to 15 times. The centrifuge does not filter well at room temperature (high concentration of immersion liquid and high viscosity), and filtration at 70 ° C can prevent crystallization. The immersion liquid is sent to the electrolytic recovery lead and bismuth, and the crystallization of arsenic and tin is recovered, and the intermediate product is sent for separation and purification. The alkali leaching residue is washed and filtered, and then sent to reduction smelting. In the smelting, the slag has good fluidity and can produce about 25% of silver containing lead.

Seven, glycerol alkali leaching method

The method has carried out glycerol-alkali leaching tests on lead anode mud and lead-copper mixed anode mud. The test results are as follows: the ratio of acid to copper and lead anode mud is 1:10, and the storage of three kinds of anode mud is compared for six months and three months. The anode mud is leached in an alkali solution of glycerin 200g/L, NaOH 100g/L and a temperature of 85°C for 2h to dissolve the ruthenium metal, and the precious metal is left in the slag, and the solution is replaced by lead powder, and the lead product is obtained by electrowinning. The lead-bismuth alloy is arranged; after the gold-silver-rich slag is washed, gold and silver are further prepared; the washing liquid is concentrated and cooled to produce sodium arsenate. The leaching rate of base metal: Pb 88.1%, As 96.5%, Bi 87.2%, Cu 26.35%. The recoveries of gold and silver and other valuable metals were: Au99.68%, Ag99.79%, Pb84.6%, Bi87.9%. The glycerol alkali leaching process does not corrode equipment, and the metal recovery rate is high and there is no smoke. However, the consumption of glycerol is relatively large (3t/t silver), which hinders the popularization and application of the method.